diseño en macizos pobres
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DESIGN IN WEAK ROCK MASSES:
NEVADA UNDERGROUND MINING OPERATIONS
Tom Brady,
Spokane Research Laboratory, National Institute for Occupational Safety and Health,
Spokane, WA USA
Rimas Pakalnis
University of British Columbia, Vancouver, BC, Canada
Lyndon Clark
Barrick Goldstrike, Elko, NV USA
Abstract
A major focus of ground control research presently being
conducted by the Spokane Research Laboratory of theNational Institute for Occupational Safety and Health
(NIOSH) is to incorporate data on weak rock masses into
existing design relationships, with an emphasis on updatingthe span design curve for manned entries and the overbreakcurve for longhole entries. Both curves were originally
developed at the University of British Columbia, Vancouver,
BC. The original database has been augmented by informationfrom mines throughout the United States, Canada, Australia,
and Europe. The common factor in all these mines is the
presence of a weak back and/or walls. In most cases, the ore
zone is the weakest rock unit and must be stabilized so thatthe mineral-bearing rock can be extracted safety. The current
NIOSH research attempts to provide rock mechanics tools to
assist a mine operator in making economic decisions that will
also ensure a safe working environment. This paper
documents the Nevada database with a special emphasis onNevada underground gold mines.
Introduction
Many of the underground gold mines in Nevada arefound in very weak ground that creates difficult and
hazardous mining conditions. A comparative analysis by the
Mine Safety and Health Administration (MSHA) for the years1990 through 2004 (figure 1) shows that the number of
injuries from roof falls in 13 Nevada underground gold mines
has varied from a low of eight in 1990 to a high of 28 in both
1995 and 1997 (3). This high injury rate was the prime
motivator for initiation of the present study by NIOSH. Thegoal is to address the extremely difficult ground conditions
associated with mining in a weak rock mass and provide mine
operators with a database that could lead to a betterunderstanding of the mechanics of these conditions.
Figure 1.Injuries from rockfalls in Nevada underground miningoperations (after MSHA [3])
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Figure 2.Schematic showing jointed rock masses associated withNevada
Figure 2 shows schematically that existing databases arebased largely on stronger, less-fractured or altered rock
masses than those found in Nevada mines. While design
methods for excavating in a weak rock mass do exist (2), they
are geared to caving operations.
The primary mining method used in Nevada is a form ofmechanized cut-and-fill. This method is amenable to
achieving a high rate of recovery when mining irregular oregeometries. Variations on this method are practiced at
different mines and within different areas of the sameoperation, depending upon rock quality. Underhand cut-and-
fill mining (3) is employed and larger spans are required
under a weak back where the costs of installing roof supportwould be high. For purposes of this study, a rock mass is
considered to be weak if it has a rock mass rating (RMR) (4)
of less than 45% and/or a rock mass quality rating (Q) under
1.0.
In areas where rock quality allows, the mines employ aversion of longhole open stoping (5). Where spans from the
footwall to the hanging wall exceed 7.6 m, transverse open
stoping is employed with the use of primaries and secondariesadjacent to previously cemented rockfill panels.
Underground mining methods as practiced in Nevada
dictated which specific databases and stope design curvesNIOSH would focus upon. Rock mass values were calculated
during mine visits and varied from an RMR high of 70% and
a low of 16% in gold-bearing fault gouge.
Several rock mass design curves developed by the rock
mechanics group at the University of British Columbia (6) areavailable, but they were not thought to be relevant to the
mining methods employed within the weak ground of Nevada
gold mines and therefore were not updated for weak ground.
Research commenced with visits to Nevada operators inJune 1999 to address concerns and determine where NIOSH
would be able to assist. The first technical site visit was onJune 12, 2002, and initial data were collected (figure 1). Themajor objectives were to obtain information on weak rock
masses and incorporate this information into existing design
curves (7) for back spans of manned entries and a stability
graph (8) for longhole wall design.
The distribution of the original databases was based onCanadian mining data, as summarized in figure 3, and shows
how few data exist for weaker rock masses.
Span Design, Man Entry
The initial span curve was developed by the
geomechanics group at the University of British Columbia to
evaluate back stability in cut-and-fill mines. It consists of twostraight lines that divide a graph into three zones: stable,
potentially unstable, and unstable. The database for this graph
initially consisted of 172 data points from the Detour Lake
Mine of Placer Dome, Inc., in Ontario, with most of the pointshaving RMR values in excess of 60% (7). The database was
expanded to 292 observations in the year 2000 with case
histories from an additional six mines (9).
The successful use of empirical design techniques isbased upon interpolation rather than extrapolation. Thus a
decision was made to develop a database for a critical span
curve in weak rock masses. The term critical span refers to
Figure 3. Distribution of original database for back span design(top) (7) and stability graph (bottom) (8)
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Figure 4.Critical span curve adjusted to Nevada operations
the largest circle that can be drawn within the boundaries of
the excavation when seen in plan view (figure 4). The term
design span refers to spans that have no support and/or
spans incorporating a limited amount of local support (forexample, pattern bolting in which 1.8-m-long mechanical
bolts are installed on a on a 1.2- by 1.2-m pattern). Local
support is deemed as support used to confine blocks that may
be loose or that might open or fall because of subsequent min-ing in surrounding areas. The new database included RMR
values from 24% to 87% with 63% of the cases over 60%.
Less than 10% of the RMR values fell below 45%, and less
than 20% fell below 55% (10). An additional 44 observationswere added to the critical span curve from Nevada operations
(figure 4; table 1), of which 35 had an RMR less than 45%.
A brief description of the use of the critical span curve is
presented; however, the reader is referred to the detailed
reference as outlined by Pakalnis (6).Excavation stability is classified into three categories;
each category is further divided into three subcategories.
1. Stable excavation (S)
a. No uncontrolled falls of ground have occurred.
b. No movement of the back has been observed.c. No extraordinary support measures have been
employed.
2. Potentially unstable excavation.
a. Extra ground support has been installed to prevent
falls of ground.b. Movement has occurred in the back.
c. Increased frequency of ground movement has been
observed.3. Unstable excavation (U)
a. Area has collapsed.
b. Depth of failure of the back is 0.5 times the span (in
the absence of major structures). Within a weakrock mass, the depth of failure has been noted as one
times the span and sometimes even greater.
c. Limited local support was not effective in
maintaining stability.
Table 1.Nevada mines database: Back spans in weak rock
RMR (%) Span (m) Condition Other
Mine 1
45 5.5 S Stable with support
45 9.0 S Stable with support
40 6.0 S Stable with supportMine 2
40 4.0 S Stable with support
45 4.3 U Caved with support
30 3.7 S Stable with support
Mine 3
40 7.0 S Stable with support
45 2.1 S Stable with support
26 2.1 S Stable with support
25 4.6 U Caved with support
55 7.6 S Stable with support
45 3.0 S Stable with support
Mine 4
70 4.6 S Stable with support
40 4.6 S Stable with support25 4.6 S Stable with support
55 5.5 S Stable with support
30 6.1 S Caved upon longhole
30 6.1 S Caved upon longhole
45 4.6 S Stable with support
50 6.1 S Stable with support
70 11.3 S Stable with support
25 7.3 S Stable with support
30 3.0 U Prior to support placement
30 1.8 S Prior to support placement
50 6.1 S Stable with support
55 7.6 S Stable with support
55 6.1 S Stale with support
Mine 5
30 3.0 S Stable with support
30 4.3 U Caved with support
20 5.8 U Caved with support
15 3.7 S Stable with support
Mine 6
45 4.3 S Stable with support
40 6.1 U Caved with support
40 4.9 S Stable with support
Mine 7
40 4.6 S Stable with support
35 4.6 S Stable with support
Mine 8
25 5.0 U Caved. Had to spile
20 1.2 S No. support. Maximum
round possible25 2.4 S No support. Maximum
round possible
35 3.1 S No support. Maximumround possible
55 3.7 S No support. Typical round.No problems
35 4.6 S No support. Typical round.No spile/shotcrete
20 7.6 U Caved. Had to spile.
45 6.0 S Stable with Split-Sets only
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A minus-10 correction factor is applied to the final RMR
when evaluating rock with shallow-dipping or flat joints.However, the applicability of this factor in weak ground is
being reassessed because of its amorphous nature. Where
discrete ground wedges have been identified, they must be
supported prior to employing the critical span curve. Stabilityis generally defined in terms of short-term stability becausethe database is based largely on stoping methods that, by their
nature, are of short duration. Movement of the back greater
than 1 mm within a 24-hour period has also been defined as a
critical amount of movement for safe access (6). This value isalso being addressed for weak rock masses as it applies to the
initial database identified in figure 2. This critical value maybe much greater than 1 mm.
Stability Graph MethodNonentry
The original stability method for open stope design was
based largely on Canadian operations and was proposed in
1981 by Mathews (11), modified in 1988 by Potvin (12), andupdated in 1992 by Nickson (13). In all instances, stability
was qualitatively assessed as either being stable, potentially
unstable, or caved. Recent research at the University of
British Columbia has augmented the stability graph by using
stope surveys in which cavity monitoring systems were em-ployed (8). This research has enabled the amount of dilution
to be quantified. A parameter termed the "equivalent linear
overbreak/slough" (ELOS) was introduced by Clark (8) andwas used to express volumetric measurements of overbreak as
an average depth over an entire stope surface. This has
resulted in a design curve as shown in figure 5.A limited number of observations existed for RMR
values under 45% (figure 3-bottom). An additional 45 data
points were added on the stability graphnonentry from
Nevada operations having an RMR under 45% (figure 6; table
2). In addition, mine 4 reflects over 338 observations thathave been averaged to reflect the design points shown in table
2. The stability graph relates hydraulic radius of the stope wall
to empirical estimates of overbreak slough. Hydraulic radiusis defined as the surface area of an opening divided by
perimeter of the exposed wall being analyzed.
The following equation was employed for calculation of
parameters for the database shown in figure 5.
N = Q * A * B * C
where N = modified stability number,
Figure 5.Stability graph (after 8)
Q = modified NGI rock quality index (14)where the stress reduction factor and
joint water reduction factor are equal to
1, as they are accounted for separately
within the analysis,A = stress factor equal to 1.0 due to relaxed
hanging wall,
B = rock defect factor. This value results
from parallel jointing and the amorphousstate of the weak rock mass being set to
0.2 and 0.3, respectively, as in table 3.
and C = stope orientation factor as defined infigure 5, that is, C = 8 - 6 cos (dip of
hanging wall).
Figure 6.Wall stability graph as developed for Nevada operations
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Table 2.Nevada mines database: Stope spans and walls in weak rock. All values of A = 1.
RMR (%) Dimensions, height bylength (m)
Dip B C N HR (m) ELOS (m) Comments
Mine 1:
45 20 by 17 90 0.3 8 2.7 4.6
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Table 3.Nevada mines database: Underhand cut-and-fill
Mine Amount of cement(%)
Span (m) Sill thickness(m)
Unconfined compressivestrength (MPa)
Comments
1 10.0 6.1 3 2.0 Paste
2a 6.5 7.6 4.6 5.5 Cemented rockfill
2b 8.0 9.1 4.6 6.9 Cemented rockfill design
2c 8.0 21.0 4.6 6.9 Mined remotely. No cave.
3 70 3.0 3. 4-12 Cemented rockfill
4a 9.0 13.7 4.0 8.3 Cemented rockfill testpanel
4b 9.0 3.7 3.0 8.3 Cemented rockfill drift andfill
4c 9.0 7.3 3.0 8.3 Cemented rockfill panel
5 *7.0 2.7 3.0 3.4 Cemented rockfill
6 6.75 4.9 4.3 3.4 Cemented rockfill
7a 10.0 1.8 2.7 0.3 Paste (factor of safety =1.5)
7b 2.4 2.7 0.57c 3.0 2.7 0.7
7d 3.7 2.7 1.0
7e 4.3 2.7 1.4 7-day sample
7f 4.9 2.7 1.8
7g 5.5 2.7 2.3
7h 6.1 2.7 2.9
8 7.0 4.6-6.1 4.6 5.5 Cemented rockfill
9 10.0 5.0 5.0 4.45 Cemented rockfill
10 12.8 6-9 6.0 2.0 High-density slurry (78%solids by weight)
11 10.0 3.0 3 (includes 0.9-m air gap)
2.5 (after 7 days) 10% cemented hydraulicfill (73%-75% solids byweight)
12 8.0 2.4-4.6 3 (includes 0.6-m air gap)
4.8 8% paste (no free water)
* Includes fly ash binder.
An initial observation from figure 6 is that the classical
design curves (ELOS) as shown in figure 5 are inaccurate at
low N' and hydraulic radius values. If hydraulic radius is keptbelow 3.5 m in a weak rock mass, the ELOS value should
remain under 1 m. It appears a hydraulic radius under 3 m
would not result in ELOS values much greater than 1 m. Thisresult is being further evaluated.
Underhand Mine Design
As noted earlier, most mines use a form of cut-and-fill
mining for a major portion of their production. As groundconditions become weaker, the primary method is mechanized
underhand cut-and-fill. These methods assure a high degree of
gold recovery under an engineered back of cemented rockfill
and/or cemented paste fill (3). To achieve the most production
at the lowest cost while maintaining a safe work environment,
mine personnel were interested in developing support
methods for the back that would best utilize either cementedrockfill or cemented paste fill. The result was the development
of a new empirical database of successful underhand mining
scenarios for the weak ground of Nevada.A major study is also underway at the University of
British Columbia to develop design guidelines for mining
under paste backfill. Figure 7 is adapted from Stone (15) and
is largely based on fixed-beam bending as the critical failuremechanism. The guidelines relate material properties to the
thickness of the beam, the exposed span, and the resultant
unconfined compressive strength required. A factor of safety
in excess of 2.0 is incorporated. The analytical relationship
was augmented by underhand cut-and-fill and shows thatactual mine design parameters generally exceed those
required by simple beam theory. The database is shown in
figure 7 and summarized in table 3.
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Figure 7.Underhand mine design curve relating span, sill thickness,and fill strength
Figure 8.Structurally controlled wedge
Support Capacity Guidelines
The development of support capacity guidelines is criticalto the overall success of the mining method selected in terms
of ensuring a safe work place (see figure 2). Ground support
in weak rock presents special challenges. Underdesign can
lead to costly failures, whereas overdesign can lead to highcosts for unneeded ground support. Figure 8 depicts a classic
wedge failure controlled by structure. It is critical to design
for the dead weight of the wedge in terms of the breaking load
of the support, as well as the bond strength associated with
embedment length (10).Over 400,000 Split-Set (10) friction bolts are used in
Nevada mines as primary support. Friction bolts are
particularly useful in fissile, buckling, or sheared groundwhere it is difficult to secure a point anchor. Caution must be
used when using this method of primary support because of
the low bond strength between broken rock and the bolt andbecause of the susceptibility of the bolt to corrosion. In mine
4, Split-Set bolts had a life of 6 months because of corrosion
resulting from acidic ground conditions. An analysis of the
performance of friction bolts in mines with weak rock (as
determined by RMR) needed to be addressed. With one
exception, Nevada mines use 39-mm Split-Set bolts (theexception uses 46-mm Split-Set bolts); mines in Canada,
however, use 33-mm Split-Set bolts. Canadian minesgenerally use these bolts only in the walls and not in the back.
Table 4 shows an updated support capacity chart as
augmented by this study.
Data points gathered from several pull tests in weak rockwere plotted as shown in figure 9. The graph shows a strong
trend between RMR and bond strength; this relationship is
being assessed as part of on-going research. Preliminaryresults are shown in figure 10.
Variability in test results shows the difficulty in assessing
overall support for a given heading. Thus, it is important that
mines develop a database with respect to the support used sothey can design for variable ground conditions. Factors
critical to design, such as bond strength, hole size, support
type, bond length, and RMR, should be recorded so as to
enable where they lie on the design curve to be determined.
Conclusions
The Spokane Research Laboratory and the University ofBritish Columbia geomechanics group are focusing on the
development of safe and cost-effective underground design
guidelines for weak rock masses having an RMR in the rangeof 15% to 45%. Weak ground conditions, ground support, and
mining methods used in several Nevada underground mines
were observed. The RMR values were calculated to updateboth span design calculations and stability graphs, and a data-
base on underhand mining methods was developed to reflect
existing Nevada mining conditions. The immediate rock mass
was also characterized and analyzed in terms of prevailing
type of ground support, potential failure mechanisms, and
rock behavior.Variability in field conditions show the difficulty in
assessing overall support for a given heading. It is imperativethat mines develop their own databases based on the type of
support used in their mines so unexpected ground conditions
can be analyzed.
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Table 4.Nevada mines database: Back spans in weak rock
Rock properties, tonnes Screen Bag strength, tonnes
Bolt strength Yield strength Breaking strength 4- by 4-in welded mesh, 4 gauge 3.6
5/8-in mechanical 6.1 10.2 4- by 4-in welded mesh, 6 gauge 3.3
Split-Set (SS 33) 8.5 10.6 4- by 4-in welded mesh, 9 gauge 1.9
Split Set (SS 39) 12.7 14.0 4- by 2-in welded mesh, 12 gauge 1.4
Standard Swellex NA 11.0 2-in chain link, 11 gauge, bare metal 2.9
Yielding Swellex NA 9.5 2-in chain link, 11 gauge, galvanized 1.7
Super Swellex NA 22.o 2-in chain link, 9 gauge, bare metal 3.7
*20-mm rebar, No. 6 12.4 18.5 2-in chain link, 9 gauge, galvanized 3.2
*22-mm rebar, No. 7 16.o 23
*25-mm rebar, No. 8 20.5 30.8
No. 6 Dywidag 11.9 18.0
No. 7 Dywidag 16.3 24.5
Note: 4 gauge = 0.23-in diameter; 6 gauge = 0.20-in diameter;9 gauge = 0.16-in diameter; 11 gauge = 0.125-in diameter; 12gauge = 0.11-in diameter
No. 8 Dywidag 21.5 32.3 Shotcrete shear strength = 2 MPa (200 t/m2)
No. 9 Dywidag 27.2 40.9 Bond strength
No. 10 Dywidag 34.6 52.0 Split-Set, hard rock 0.75-1.5 mt per 0.3 m1/2-in cable bolt 15.9 18.8 Split-Set, weak ground 0.25-1.2 mt per 0.3 m
5/8-in cable bolt 21.6 25.5 Swellex, hard rock 2.70-4.6 mt per 0.3 m
1/4 by 4-in strap 25.o 39.0 Swellex, weak rock 3-3.5 mt per 0.3 m
Super Swellex, weak rock >4 mt per 0.3 mNote: No. 6 gauge = 6/8-in diameter.; No. 7 gauge = 7/8-in diameter.; No. 8 gauge = 1-in diameter. 5/8-in cable bolt, hard rock 26 mt per 1 m
NA = Not applicable. No. 6 rebar, hard rock 18 mt per 0.3 m, ~12-in
granite
Figure 9. Pull-out load versus RMR for SS39.
The results from augmented design curves and pull-outtests are presented in the hope that they will aid mine
professionals in their task of designing a safe workplace. A
systematic approach allows an operator to understand overall
failure mechanisms and resultant loads that could affect thesystem. This approach would allow an engineer to develop an
optimal support strategy for the mining method employed.
The work would not have been possible without the
partnership between NIOSH, the University of British Colum-
bia geomechanics group, and Nevada gold mining company
personnel. This continued partnership is critical to thedevelopment of safe and cost-effective mine strategies. Figure1 shows that since the inception of the team approach and
resultant collaboration, injury statistics have declined dramati-
cally. This decline may be a result of many factors; however,
it is clear that this approach is important and relevant to mine
operations.
Figure 10. Pull-out load versus RMR at mine 4
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Acknowledgments
Specific thanks go to Patrick Carroll, chief engineer,
Newmont Midas Mine; J.J. Hunter, chief rock mechanics
engineer, Newmont East Carlin and Deep Post Mine; MonicaDodd, chief engineer, Newmont Deep Post Mine; Simon
Jackson, chief engineer, Placer Dome, Turquoise Ridge Joint
Venture; Shawn Stickler, chief ground control engineer,
Queenstake Murry Mine; and Rad Langston, senior rockmechanics engineer, Stillwater Mining Company.
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